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By Z. Zhang, M. S. Gonzalez, M. de P. Bueno, P. Holopainen, S. Luukkanen

"Despite being a field of study that requires practical knowledge and experience, mining engineering education follows the same pattern in most mining universities around the world, consisting of theoretical courses and laboratory-based practical modules (Mischo, 2015). The possibility of having practical training in real mining conditions on actual mine sites or processing plants is limited, while the only contact that undergraduate students have with mining operations during their time of study is usually through shorttime internships or visits in the context of educational trips. Real integration between universities and the mining industry may be scarce, which does not help students to observe the actual processes and thus understand the whole picture of mining operations.Indirect practical training is now possible since that technology is shifting from mechanical to digital, and the use of virtual reality — in the context of Mining 4.0 — can provide additional experience to the learning process. However, virtual reality is limited to the visual impression and is also limited in actions (Binder et al., 2018). In any case, when there is not a proper pragmatic view of how the different disciplines link to each other, especially if there is a lack of communication, unfortunately the results could be seen at a later stage in the mines as a poor performance of professionals and workers.Another major challenge for educating specialists in the mining sector is that experimental mines suitable for practical education, where the skills and knowhow could be developed and/or enhanced, are rare worldwide. Such research and educational (R&E) facilities are practically former metal ore mines that have been transformed into experimental sites. Besides being limited in number, not all facilities of this kind serve both a research and educational purpose.An early example of asolely experimental mines is the Luossavaara-Kiirunavaara Mine site operated by LKAB in Sweden, where systematic research on rock mechanics and mining engineering was successfully carried out in the 1980s. As a result, the developed underground mining technology was applied in LKAB’s mines at a later stage. There have also been research mines within collieries such as the former Lake Macquarie Mine at New South Wales in Australia and the Zeche Tremonia Mine in Germany), but these are no longer in operation (Mischo, 2015)."

Oct 1, 2019

By Xiaochao Shu, Kun Wang, Fanwei Meng, Rui Tao

Rapidly searching for carbonatite-related rare earth element (REE) mineralized systems is of paramount importance to explorers, because these systems provide the world’s most significant supply of REE. Biotite and barite are suitable indicator minerals that can offer reliable elemental geochemical information used for REE exploration. In this study, we provided an elemental geochemical dataset of biotite and barite from Cenozoic REE deposits in the eastern Tibetan Plateau, and compared the obtained data to those of biotite and barite forming in other geological settings. The F (i.e., fluorine) contents in carbonatite-hosted biotites are highly variable, with biotites from mineralized carbonatite complexes having much higher values (1.05–2.37 apfu, where apfu = atoms per formula unit) than their counterparts from barren carbonatite complexes (≤ 1 apfu). It is therefore believed that the presence of high-F biotites can be regarded as a sign for predicting REE reservoirs. Barites in REE mineralized systems contain much higher (even one or to orders of magnitude) total REE contents (ΣREE >100 μg/g) relative to those species forming in carbonate-hosted fluorite-barite deposits and open marine environments. This means that the high ΣREE concentration of barite can serve as an effective index for REE exploration. The above information is integrated into a REE exploration program in which the areas with strong local tectonism along cratonic edges are priority exploration targets, and high-F biotites and REE-rich barites can serve as direct indices for rapidly identifying carbonatiterelated REE mineralized systems at a strategic distance from other geological settings.

Jun 9, 2023

By J. S. Gardner, J. J. Zaluski

I. INTRODUCTION West Virginia has a long history of coal production.1 The state at one time led the nation in total coal production, while at the same time coal from West Virginia constituted one quarter of total U.S. coal production.2 In more recent years, however, West Virginia’s level of coal production has been reduced in comparison to that of other states due to the high relative cost of producing coal in the state.3 Another blow to the producers of coal in West Virginia is the action known as Bragg v. Robertson.4 In 1998 various individual plaintiffs, as well as a conservation group known as the West Virginia Highlands Conservancy, filed a civil action under a provision of the Surface Mining Control and Reclamation Act of 1977, Pub. L. No. 95-87, §522, 91 Stat. 445 (codified at 30 U.S.C. §§1201-1328) ("SMCRA" or the "Act") that allows declaratory and injunctive relief against state and federal officials for non-compliance with the Act.5 The plaintiffs generally claimed that the Director of the West Virginia Department of Environmental Protection ("WVDEP") had exhibited a pattern of violating mandatory non-discretionary duties under SMCRA and the West Virginia state regulatory program.6 These non-discretionary duties involved many areas of regulation, including water quality standards, disturbance of wetlands, hydrologic reclamation plans, Approximate Original Contour ("AOC") requirements and AOC variances, post-mining land uses, contemporaneous reclamation, and as explained later, the "buffer zone rule."7 In addition to these claims against WVDEP, the plaintiffs also alleged that certain individual members of the Army Corps of Engineers (the "Corps") had failed to carry out statutory duties under the Clean Water Act (the "CWA") and the National Environmental Policy Act ("NEPA").8 Generally, the complaint stated that the Corps did not have statutory authority under the CWA to regulate valley fills created for the purpose of the disposal of waste material, i.e., the overburden associated with mountain top mining.9 Alternatively, the plaintiffs claimed that even if the Corps could regulate valley fills, the Corps violated NEPA by issuing nationwide permits without required analysis. Finally, the plaintiffs asserted that the issuance of a nationwide permit for surface mining valley fills is unlawful.10

Jan 1, 2001

It is widely accepted that particles stabilise flotation froths and that stable froths result in improved flotation performance. Predicting the effect of particle addition on froth stability is, however, challenging. Dynamic surface tension measurement using maximum bubble pressure presents an attractive technique to investigate the effect of surfactant and particles at the air-water interface. The range of bubble lifetimes that can be studied (typically 0.1 to 60 s) is analogous to variations in air rate in flotation cells, and the corresponding changes in surface tension give an indication to the diffusion and adsorption rates of particles at the interface. In this paper, we use dynamic surface tension measurements to investigate the effect of particles on bubble surfaces at the microscale and link this to bulk froth stability measurements carried out using a froth column. Using the maximum bubble pressure method, the results show that the addition of particles results in lower surface tension, both at the dynamic (i.e. short) bubble lifetimes and towards equilibrium (i.e. 60 s bubble lifetime). This corresponds with the bulk froth stability measurements, where the three-phase system yielded more stable froths than the surfactant only system. Furthermore, increased particle loading at the air-water interface, whether through higher surfactant concentrations or lower air rates (longer bubble lifetimes), gave lower surface tension and higher froth stability. This demonstrates the link between bubble loading and froth stability. It is proposed that the maximum bubble pressure technique can be used to predict froth stability for two- and three-phase systems, enabling the effect of particle loading to be accounted for and quantified. Moreover, the technique has the potential to allow rapid determination of particle and surfactant diffusion at the air-water interface and prediction of the corresponding effect on bulk froth behaviour.

By M. Sugamata

Creep tests were carried out for friction welded joints of AZ31 magnesium alloy rods under an ambient atmosphere at temperatures between 423 and 573K. Tensile tests were also performed at temperature from room to 573K. The obtained results were compared with those of the base metal. AZ31 extruded rods of 19 mm in diameter were joined by brake type friction welding machine. Creep and tensile specimens were machined out of the welded rods. The tensile strength of the joints was almost the same as that of the base metal at room temperature. However, joint efficiencies of welded joints were about 92% at above 373K. The elongation of welded joints was lower than those of base metal. The welded joints showed appreciably higher minimum creep rate and shorter rupture time than the base metal at all the tested temperatures and initial creep stresses. Creep and tensile rupture of the joints always occurred at near the weld interface. Thus, friction welded joints of AZ31 alloy showed lower creep and tensile strength than the base metal at elevated temperature.

Jan 1, 2006

By Peter Faust

The foundations for the new South Viaduct Bridge in San Francisco were extremely challenging due to the requirement for very large and deep piles installed adjacent to historic structures. Drilled shafts up to 12? foot in diameter were constructed to a maximum depth of 200? feet in alluvial soils and Franciscan Bedrock using an oscillator system for the installation of the permanent steel casing. This paper provides an overview of the construction techniques, the equipment, and the challenges associated with constructing 12? foot diameter shafts on this very complex project in one of the nation?s nicest landscape setting.

Jan 1, 2011

By Abhishek Kumar, T. G. Sitharam, Surender Singh

The height of tailings pond (TP) embankments is frequently raised to accommodate the ever-increasing volume of tailings released by the mining industries. Excess pore water pressure (EPWP) may develop during the height-raising operations, which can affect the overall stability of TP. This article presents a detailed stability analysis for the embankments of an existing TP, the height of which was increased twice utilizing downstream (D/S) and upstream (U/S) construction techniques. A rigorous two-dimensional transient fully coupled-stress-pore pressure analysis is performed in the finite element-based package RS2 to examine the build-up of EPWP during different stages of construction. Furthermore, an attempt is made to investigate the feasibility of future height raising, with due consideration given to the influence of embankment raising rates. Results from the analysis show a development of higher EPWP within the tailings deposit in comparison to the embankments of TP, indicating the development of an undrained-like condition. The study further reveals that the development of EPWP is highly affected by the rate at which embankments of TP are being raised. A higher rate of embankment-raising results in the development of greater EPWP in comparison to a lower embankment-raising rate, which may lead to the ultimate collapse of TP. Furthermore, based on the stability analysis, the height of existing TP can be further raised by 4 m (two lifts of 2 m height) utilizing the U/S construction technique. A thorough cost and benefit analyses are however required before implementing the results of the current analysis since the storage capacity of TP is reduced while raising by the U/S method.

May 16, 2023

By J. J. Harrington

The pH control system at Kennecott Copper Corporation's Nevada Mines Division concentrator consists of two independent systems, lime dilution and lime addition. Lime dilution is accomplished in a mixing tank where slaked lime and water are introduced proportionately. The desired dilution of the milk of lime is determined by the electrical inputs provided by eight pH probes, one per grinding section. The inputs are averaged and applied to a time base generator. The summation output of the time base generator determines the proportion of lime and water that are to be mixed and applies this input to a voltage sensitive switching device. The switching device provides a signal to one of two relays which will add either slaked lime or water, dependent upon the summation output. The entire cycle is based on a two-minute repetition rate to allow for sufficient mix time. Responsivity is controlled by maintaining a limited maxi- mum solution level in the mix tank. Lime addition is controlled independently within each grinding section. The electrical input from the pH probe in a given section controls the solenoid valves which provide lime to each of the ball mills. The valves remain open within a time range capable of maintaining the pH of flotation feed at a given set point.

Jan 1, 1975

By D. Imeson, T. Grammatikopoulos, G. Pearse, M. Aghamirian, C. Mohns

"Spodumene mineral is a source of lithium for lithium ion batteries. Thus, the production of a spodumene concentrate with an approximate grade of 6% Li2O is targeted. Traditionally, spodumene is concentrated by froth flotation using fatty acid as the collector. Amine collector can also be used to float mica ahead of spodumene as required or feldspar from the spodumene tailings. Slimes have a negative effect on spodumene flotation and they are commonly separated by scrubbing and de-sliming prior to spodumene flotation. The main losses in spodumene flotation typically occur in cyclone overflow, scrubbing slimes, spodumene flotation tailings and mica concentrate. In recent years, dense media separation (DMS) is the main concentrating process in Australian spodumene operations. In these operations, DMS is efficient at generating coarse grain concentrates and rejecting large portions of silicate gangue, thereby directly reducing the milling cost. The ore should be crushed to 3/8” or coarser while further crushing is possibly necessary on the middling particles to scalp further spodumene. The DMS feed should be classified to remove fines (typically -0.5 mm particles). The final middlings and undersize fractions could then be processed by flotation. The main requirement in DMS operations is to have good spodumene liberation at relatively large grain size. This paper reviews the most significant DMS and flotation parameters in spodumene beneficiation. The treatments of different minerals in the spodumene ore are also addressed.INTRODUCTION A number of laboratory/pilot plant programs have been recently conducted at the SGS Lakefield site, sponsored by different clients. The main objective has been to develop a process flowsheet to generate a spodumene concentrate containing 6.0% or higher Li2O (~75% spodumene) suitable for producing lithium carbonate by a hydrometallurgical process. Depending on spodumene liberation size, dense media separation (DMS) can be an attractive option to produce a final high grade concentrate and to reduce the mass to be treated in the flotation circuit. In such a case, crushing would typically be performed in stages, with DMS separation to recover a spodumene concentrate and reject gangue minerals following each stage. DMS middlings and -0.5 mm fines can be processed by flotation to further recover spodumene. In the flotation process, coarse stage grinding minimizes the grinding cost and formation of slimes. In the majority of cases, a grind size (K80) of approximately 200 microns is sufficient for flotation. After primary grinding, mica, if present in an appreciable amount, should be floated to minimize interference with spodumene flotation. Desliming after scrubbing and high density conditioning are the key process operations prior to spodumene flotation. Spodumene can then be floated from silicate gangue minerals with a fatty acid or a mixture of fatty acid and petroleum sulfonate at slightly alkaline pH. Subsequently, feldspars can then be floated from the spodumene tailings."

Jan 1, 2012

By S. M. Cohen

Roll crushers had been used in the past ahead of existing milling systems in order to obtain additional production from such mills in both the mining and the cement industries. Historically, a roll crusher as a preliminator to a ball mill has been used for a number of industrial mineral applications. Flowsheets, from Taggars(1) show systems for fluorspar, graphite, lead and zinc to name a few, which are shown in Figures 1 to 4. These units applied, (in some cases still in service), prior to 1950 and as early as the 1920's, were of two general types, rigid rolls and spring rolls. The rigid rolls were phased out in favor of the spring rolls. The spring force related to the linear inch width of the rolls, ranged from 500 to 50,000 lbs/linear inch(1) on units up to 90" diameter. With choke feeding, the rolls were spread against the spring force during crushing keeping the system under high pressure to obtain high effectiveness.

Jan 1, 1986

By Patterson B. G

The small party of mining men who formed the original Mount Morgan Syndicate in 1882, commenced operations on the top of the mountain with what were practically nothing more than small hillside quarries, It was not until 1891 that underground stoping began, and open-cut mining therefore commenced with the commencement of operations. This hillside quarrying, however, seems to have been quite haphazard and unsystematic ; there is no trace or sign of any properly devised mining scheme. This can be seen plainly enough from an examination of old mine plans; for in less than five years after stoping was; adopted, the open-cut was down into the stoped-out ground. The following table, prepared from these old plans, shows the downward trend of the excavation in the past fiftyseven years:-

Jan 1, 1939

By D. C. Shumway

Mining in Cushenbury Canyon began in 1860 when John Cushenbury left his Holcomb Valley gold prospect and discovered a silver vein in the limestones of the present day Cushenbury Mine. Cushenbury Canyon is located along the western boundary of the Mojave Desert in Southern California and on the north slope of the San Bernardino Mountains. The Holcomb Valley, immediately southeast of Cushenbury Canyon was the most prolific gold producer in Southern California. Gold mining dominated the area until World War II when interest in the Lucerne Valley Limestone deposits was peaked by the construction of the Kaiser Steel Plant in Fontana, California. The steel making process required large amounts of metallurgical grade limestone and Kaiser geologists located claims in most of the limestone deposits. Kaiser Cement constructed a cement plant and limestone mine in Cushenbury Canyon in 1956. Kaiser Cement modernized the facility in 1982 to a four stage preheater-precalciner system. Mitsubishi Cement Corp. purchased the property in April, 1988, and through further system modifications have created the most efficient and modern cement plant in the United States.

Jan 1, 1992

By James L. Coburn

Leaching is defined as a form of extraction in which a valuable constituent is removed from a solid body by mass transfer to a solution which, for the purpose of this discussion, is considered to be an aqueous medium. Historically, leaching has been applied to the recovery and purification of many elements. This paper, however, is directed toward the recovery of copper, uranium, vanadium, and precious metals included in the following systems: 1. The leaching of copper minerals, sulfides and oxides, by acid or by ammoniacal leach solutions with or without an oxidant. 2. The leaching of gold and silver, either native metals or as minerals by an oxygen-containing alkaline cyanide solution. 3. The solubilization of certain uranium and vanadium minerals by oxygen or oxidant-bearing lixiviants. Many other systems of this general type exist, but from a production standpoint, these three systems account for most of the major values recovered by leaching.

Jan 1, 1975

By A Kondratiev, A E. Moiseev

Viscosity of oxide melts is one of the fundamental physicochemical properties that plays a crucial role in important technological and natural processes like slag flow, slag/metal separation, volcano eruptions etc. However, many oxides melt at extremely high temperatures and are highly corrosive in the liquid state, which makes experimental measurement of melt viscosity by traditional experimental techniques a difficult and expensive, if not unachievable, task. In this case it might be useful to employ other methods of viscosity determination such as modelling or simulation. In the present paper the shear viscosity coefficients of selected pure refractory oxides (CaO, MgO, Al2O3 and others) have been simulated at temperatures above their melting points via the classical molecular dynamics and compared to the available viscosity data (eg experimental viscosities, other model predictions) collected by the authors. The simulation has been carried out using the noncommercial molecular dynamics simulation package LAMMPS and the Born-Mayer-Huggins potential, which parameters for the selected oxides were taken from the literature. Several viscosity calculation techniques (the Stokes-Einstein and Green-Kubo equations, the so called Einstein relations, and the non-equilibrium molecular dynamics methods) have been applied to ensure a more reliable viscosity calculation. It has been demonstrated that the simulated viscosity of the selected melts agrees with the available experimental data. It has also been shown that the simulated viscosity of a number of the oxide melts investigated is close to that calculated by phenomenological viscosity models, which supports viscosity extrapolation from the corresponding binaries and ternaries, if no experimental data is present for unary systems. In general, it has been demonstrated that the molecular dynamics simulation could provide a reasonable estimation of viscosity if no other viscosity data is available.

Jun 19, 2024

By George Breyer

DURING the past 25 years the steelmaker has experienced some difficulty in producing steel to meet the standards established by the metallurgist. In contrast to the past, when a chemical specification of broad limits was the only requirement, he now has to employ such tests as microscopic, macroscopic, tensile, impact, Magnaflux and even X-ray and supersonic. The development of these tests has more than set the pace for a manufacturing process of which the exactness still hinges to an important degree on the ability of the furnace operator. In our efforts to produce steel, and especially alloy steel, to meet the exactness of modern engineering specification, many factors must be taken into account. It is not sufficient to make a good heat of steel in the furnace and then be satisfied that the job is done, since there are many pitfalls between the furnace and the finished product. The literature abounds with information on melting practices, deoxidation methods and slag control, but the lack of positive information about what happens when the steel is teemed into molds is quite evident. Such questions as what is the proper mold size, shape and thickness; which is the best mold coating; what is the best rate of pour for any given steel, and what is the effect of cooling rates on quality, are still answered by opinions. It is acknowledged that a poor heat of steel cannot be improved in the pouring pit. On the other hand, a good heat can be ruined by improper handling after it leaves the furnace.

Jan 1, 1947

By D. B. Gawthrop

"One of the properties of an explosive, of interest to the user, is its ability to transfer detonation across an air gap to another explosive charge. This Property is of particular importance with permissible explosives used in coal mines, since they are relatively insensitive to detonation when compared to many of the commercial dynamites used in metal mining.An essential characteristic of an explosive is that it shall always be sufficiently sensitive to propagate detonation completely. An insensitive explosive constitutes a hazard because of (1) the danger of drilling into unexploded portions of charges of explosives which have partially misfired, and (2) the likelihood of a low order of detonation, or the burning, of all or part of the charge in a hole prepared for blasting. The latter condition not only usually produces large volumes of poisonous gases, but it is particularly hazardous in gassy or dusty coal mines where explosions may result.The Bureau of Mines and manufacturers of explosives determine the sensitivity of explosives to detonation over an air gap by means of a [(halved-cartridge gap test,"" for which the ends of the cartridges of explosive exposed to one another, are cut off. Since in blasting, the crimped ends remain on the cartridges used, this investigation has been made to ascertain the effect of the crimped paper ends on the distance detonation is propagated over an air gap as compared with the propagation distances obtained when the crimped-paper ends are removed."

Nov 1, 1932

By R. V. Ramani, S. Shankar

"Available literature on the effect of re-entrainment on overall airborne dust concentration in mine airways or the fundamental mechanisms of coal dust re-entrainment in mine atmosphere is scant (Shankar and Ramani, 1991 ). Therefore, as part of the Respirable Dust Center's ongoing research, the following three-part study was undertaken:• a study of the re-entrainment phenomena of anthracite and bituminous coal dust particles in a wind tunnel under the influence of varying air velocities;• a study of the same in the Lake Lynn Laboratory Mine of the USBM with both air flow and walking by personnel causing re-entrainment; and• the mathematical modeling and computer simulation of the re-entrainment phenomenon.In this paper, results from the re-entrainment studies in the wind tunnel and the Lake Lynn Laboratory mine are presented. Comparing the results from the two studies,it was observed that increasing the air velocity increased the rate of re-entrainment ( Shankar and Ramani, I 994a) and the rate of re-entrainment decreased with the time of exposure of the source dust to the flowing air. In addition, when the source dust was composed of particles of different sizes, the airborne dust concentrations due to re-entrainment were higher than when the source dust was composed of particles of the same size. The contribution of airborne dust concentration due to air flow was negligible compared to the contribution due to walking. IntroductionIn underground coal mines, dust is produced primarily during the process of mining and during such activities as pillar spall, loading, belt transfer and roof fall. The generated dust is entrained and then collects on surfaces through dispersion, diffusion and eventual deposition. Removal forces, such as flowing ventilation air or turbulence caused by walking and moving vehicles, can re-entrain the deposited dust (Hall, 1955). A portion of the re-entrained dust remains airborne and increases the concentration of airborne dust in mine airways.The presence of airborne coal dust in mine airways in large concentrations can be hazardous. Health effects include coal workers' pneumoconiosis (CWP) and silicosis, and it can be a nuisance affecting visibility. In addition, as is common during methane explosions, if the concentration of airborne coal dust becomes significant, there could be a coal-dust explosion (Singer, Cook and Grumer, 1972; Singer, Harris and Grumer, 1974; Edwards and Ford, 1988). Through concerned research efforts over the last decade, significantprogress has been made in reducing the airborne dust concentration in mine atmospheres. However, secondary sources, such as re-entrainment, have not been studied in-depth. It is important to understand the phenomenon of re-entrainment and its potential contribution to the airborne dust concentration in mine airways for further decreasing airborne dust hazards in mines."

Dec 1, 1996

By F. Fraas, H. H. Storch

"In the second paper 4 of this series, data were presented concerning the decomposi¬tion of polyhalite by saturated salt solutions. It was found that uncalcined polyhalite upon leaching with boiling saturated sodium chloride solutions yielded a brine in which the (K2 Mg) /304 ratio was considerably higher than that obtained from the calcined material, and for this reason appeared more suitable for the production of potassium chloride. It was found, in fact, that the leach liquors from uncalcined polyhalite deposited a mixture of KC1 and NaC1 upon evaporation and subsequent cooling, whereas those from calcined polyhalite yielded mixtures which contained not only potassium and sodium chlorides but also large quantities of glaserite, K3Na (S04)2. Hence it was decided to study the evaporation of the uncalcined poly—halite leach liquor and postpone experiments on the calcined polyhalite liquors until such time as a process for further removal of sulphate was developed.A considerable body of equilibrium data is available in the literature on thesystem K+ Mg++ — Na+ Cl—S04 =,— H20 for temperatures up to 83º C. Owing, however, tothe lack of data other than that corresponding to the invariant points and the complete absence of data for temperatures of 100 to 115°C., and owing to the fact that supersaturation phenomena render unsafe the prediction of the behavior of such a system upon evaporation, it was considered essential to obtain the necessary data by actual evaporation and analysis of the brines thus obtained."

Feb 1, 1931

By W. Wanjokow, H. O. Hofman

In the metallurgical treatment of most metallic sulphides it is usually necessary to carry on a roasting-operation. In some cases the raw ore will have to be roasted; in others the intermediary product, matte, produced in smelting the ore; again, both ore and matte may have to be subjected to oxidation at elevated temperatures. Most metallic sulphides, when given an oxidizing-roast, pass through the state of sulphate before they are converted into oxides. All metallic sulphates are completely decomposed by heat, with the exception of those of lead and bismuth, but the temperatures at which decompositions take place vary with the metals. Some metallic sulphates are converted directly into oxides, others first form basic salts. Alkali-earth sulphates show a behavior similar to that of metallic sulphates, while alkali sulphates do not part with their sulphur trioxide even at elevated roasting-temperatures. The older data on the decomposition of metallic sulphates by heat are only approximations. Thus Kerl' gives the order in which metallic sulphates are decomposed with increasing temperatures as (Ag2, Fe, Cu, Zn, Ni, Co, Mn, Pb) SO,. More recent statenlents give definite temperatures and modify the formerly accepted order of decomposition. But the temperatures given by various observers do not agree. The reasons for this are to be found mainly in the method and manner of carrying out experiments. Four methods have been used: 1. Heating in an evacuated tube; 2, heating in a tube closed at one end, without air-cur-

Jan 1, 1913

By M C. Goode

The use of electrical or electronic safety systems to protect personnel, and the engineering tasks associated with the design, manufacture and maintenance of these systems are referred to using the term ‘functional safety’. In the past ten years, the use of such systems has transitioned from its traditional base, in oil and gas major hazard facilities, to being a part of most large mining and materials handling projects. In particular, functional safety is rapidly becoming a significant component of Australian iron ore projects, typically in the form of emergency stops, remote isolation, drive through interlock, robot cell entry or mobile machine anti-collision systems. While these types of systems have been in use in iron ore for a long time, functional safety standards (eg AS61508, AS62061, AS4024.1503) require more rigourous demonstration of the level of risk reduction provided by the systems.The early engineering project development phases require a number of key decisions to be made around functional safety implementation. The decisions made at this point in a project will go on to dictate the required types of risk assessment and analysis, equipment selection, sizing of substations, and training and maintenance requirements for owners. For project teams that are unfamiliar with functional safety, or inexperienced in its application, an incorrect decision or failure to understand the requirements of the chosen standard can lead to a failure to meet contract requirements, as well as major cost and schedule overruns.This paper examines the implications of functional safety on early engineering studies, focusing on the following questions: How does a designer or company determine what standard is best suited to their project? What key aspects of functional safety design need to be considered during project development? And finally, how can the impact of functional safety to the project be minimised?CITATION:Goode, M C, 2017. Functional safety impacts on study stages of major projects, in Proceedings Iron Ore 2017, pp 347–352 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 24, 2017